In standard methods of obtaining metallic lead from concentrates, the standard procedure has been to treat the lead sulfide concentrates in a blast furnace. However, the pyrometallurgical procedure possesses many disadvantages and drawbacks. Primary among these disadvantages is that the process will result in some major pollution problems such as the generation of sulfur oxide gas along with substantial fuming. The fuming carries with it possible carcinogenic compounds which will contain lead, cadmium, etc. Therefore, it is necessary to provide improved and safer methods for obtaining metals such as lead in metallic or elemental form by methods which will not contribute to pollution of the air or will be safer to operate. The aforementioned lead smelting techniques will consist of roast sintering the lead sulfide concentrate whereby a major portion of the sulfur will be removed followed by melting in a blast furnace to obtain the metallic lead.
In an effort to alleviate the pollution problem, it is necessary to develop new processes for obtaining lead which will be competitive as an alternative to the conventional smelting practices. Prior work in the hydrometallurgical field resulted in developing a non-aqueous processing route whereby lead sulfide concentrates are chlorinated at temperatures above 300.degree. C. to produce lead chloride and volatilized sulfur. However, chlorination at these elevated temperatures will promote the formation of volatile chlorides of contaminating elements such as iron, magnesium, aluminum, silicon, zinc, as well as elemental sulfur which may be present in the lead sulfide concentrate. Other hydrometallurgical processes which have been developed include the use of ferric sulfate as a leach agent. In this method, the lead sulfide is sulfated to form lead sulfate. This step is then followed by carbonation of the lead sulfate to form lead carbonate and thereafter the lead carbonate is subjected to dissolution in hydrofluosilicic acid for electrolysis to metallic lead. Yet another hydrometallurgical method which is developed for the recovery of lead is based on the use of a ferric chloride medium. This method involves a leaching step whereby the lead sulfide is converted to lead chloride and thereafter subjected to steps of solubilizing, crystallization, and electrolysis.
The prior art which exemplifies some methods of recovering lead or other metals is exemplified by U.S. Pat. No. 607,287 in which sulfide ores are chlorinated at a temperature which is high enough to cause volatilization of some metal chlorides as well as forming gaseous sulfur chlorides. However, in such a process heat does not have to be added due to the exothermic nature of the reaction, but in order to produce elemental sulfur and to minimize the conversion of impure metals, the temperature must be controlled. The chlorinated product is then leached with water which would quickly turn into a chloride leach system containing all soluble metal chlorides to a high degree. The water in this case is a scrub solution to remove the volatile chlorides and sulfur chlorides from the off-gas stream. This therefore, is not a wash process but in contradistinction is a leach process in which all metal chlorides which had been substantially converted are leached, leaving behind a residuum of sulfur and only minor amounts of unreacted sulfides. This leach or wash solution acts as the prime metal recovery step whereas, as will hereinafter be shown in greater detail, the wash step of the present invention has, for its primary aim, the removal of a small amount of unwanted metals that had been only slightly converted. Furthermore, due to the high temperature chlorination of the process of the patent it is, of necessity, a non-selective chlorination.
U.S. Pat. No. 1,346,642 describes a separation process for lead and zinc using the differential action of strong hydrochloric acid in an aqueous leach system. In this process, which is a wet chlorination process, the action is effected at a temperature of about 100.degree. C., the hydrochloric acid converting the sulfides, except zinc, to chlorides and the sulfur to gaseous hydrogen sulfide. This treatment is not nearly as selective as the dry halogenation step of the present process, the wet acid treatment producing a greater amount of soluble zinc chloride. Furthermore, the sulfur does not remain as an innocuous solid in this process. The lead chloride and unreacted zinc sulfide are then separated from the solution containing the soluble metal chloride impurities such as iron and copper. A water wash is used to remove the entrained hydrochloric acid leach liquid in the initial leach solids and is combined with the acid leach solution forming the next series of wash. Therefore, this cold water wash actually becomes the leach solution on recycle and, as hereinbefore pointed out, will contain a relatively large amount of impurities. Inasmuch as this wash step is a displacement step to remove all of the free hydrochloric acid, it is not intended to remove the metal impurities per se, but rather to remove the solution which would continue to leach the impurities.
Another patent which is drawn to a method for recovering values from a sulfide ore is U.S. Pat. No. 1,943,340. This patent concerns a two-stage roast of zinc concentrates wherein the first step relates to the sulfation of the concentrate and the second stage relates to a chlorination. Zinc is removed after each roasting stage or step, however, there is no wash step described in this patent. Likewise, U.S. Pat. No. 3,961,941 is drawn to a method of producing metallic lead and silver from the corresponding sulfides. The patent teaches a two-stage ferric leach-brine leach process for lead concentrates and is a wet ferric chloride conversion of lead sulfide to lead chloride. The only wash step which is described in this patent involves the use of said step to wash chloride out of the process tailings before said tailings go to disposal. In contradistinction to the process of the present invention, the wash step has no effect on the resultant purity of the lead chloride precipitate.
As will hereinafter be shown in greater detail, especially in the examples appended at the end of the specification, it has now been discovered that metallic lead in a relatively pure state, may be produced in a simple and economical manner by a series of steps including the halogenation of a lead sulfide concentrate which has been subjected to a drying step prior to halogenation thereof, at a relatively low temperature in contrast to prior art methods which have involved chlorination steps at relatively high temperatures or in a wet system with the attendant formation of undesirable compounds.
This invention relates to a hydrometallurgical process for the recovery of metallic lead. More specifically, the invention is concerned with an improved process for obtaining metallic lead from lead sulfide sources or concentrates whereby unwanted side reactions are minimized, the impurities which are present in the concentrate being more totally unreacted or converted are removed in their original state thereby leading to the production of a lead halide in a purer form than has heretofore been obtainable, the end result being that the metallic lead which is recovered in the last stage of the process will be in a relatively purer state than has heretofore been possible.
It is therefore an object of this invention to provide an improved process for the production of metallic lead.
A further object of this invention is to provide a hydrometallurgical process for the production of relatively pure metallic lead from lead sulfide concentrates.
In one aspect an embodiment of this invention resides in a process for the production of metallic lead which comprises the steps of drying a lead sulfide source containing at least one metal impurity selected from the group consisting of iron, copper, zinc and cadmium, halogenating the dried lead sulfide source at a temperature in the range of from about 90.degree. to about 120.degree. C., water washing the halogenated mixture to remove solution halides of the slightly converted metal impurity, leaching the halogenated mixture with brine, filtering the resulting brine solution to separate elemental sulfur and residue from soluble lead halide, crystallizing said lead halide and recovering metallic lead by electrolysis.
A specific embodiment of this invention is found in a process for the production of metallic lead which comprises drying lead sulfide at a temperature in the range of from about 100.degree. to about 150.degree. C., chlorinating the dried lead sulfide by treatment with chlorine gas at a temperature in the range of from about 90.degree. to about 120.degree. C., water washing the chlorinated mixture to remove the small amount of soluble metal chloride impurities, leaching the solid residue at a temperature in the range of from about 80.degree. to about 120.degree. C. with a sodium chloride solution, maintaining the pH of the leaching solution in a range of from about 4 to about 8 by the addition of an acidic or caustic solution, filtering the leached mixture at a temperature in the range of from about 80.degree. to about 120.degree. C., crystallizing the soluble lead chloride, and recovering metallic lead by subjecting the crystallized lead chloride to a fused salt electrolysis.
Other objects and embodiments will be found in the following further detailed description of the present invention.
As hereinbefore set forth the present invention is concerned with an improvement in a hydrometallurgical process for the production of metallic lead. The feedstock which is utilized in the present process will comprise a lead sulfide source either in the form of flotation concentrates or raw feed ores which are materially rich in lead sulfide, although it is contemplated that some of the lead may be present in the form of lead carbonate or lead oxide.
In the first step of the process the feedstock is subjected to a drying operation in order to remove any water which may be present, in order that the material to be processed remains fluid during the processing operation and does not cake, and also that the evolution of water will not occur during the subsequent halogenation step to such an extent which is great enough to form significant quantities of hydrogen halide such as hydrogen chloride, hydrogen bromide, etc., or other detrimental reagents or products which could effect either the chemical or physical parameters of the process. The drying of the feedstock is effected at elevated temperatures ranging from about 100.degree. to about 150.degree. C., for a period of time sufficient to reduce the water content of the feed to a value of 2% or less. This drying of the feedstock within the temperature hereinbefore set forth differs from the prior art as exemplified by U.S. Pat. No. 3,961,941 in which the feedstock is subjected to an initial roast at elevated temperatures ranging from about 525.degree. to about 900.degree. C. under nonoxidizing conditions.
Upon completion of the drying of the feedstock and reduction of the water content the dried feed is then subjected to halogenation. In contradistinction to prior art methods hereinbefore discussed, the present invention utilizes a halogenation temperature of the lead sulfide at relatively low values ranging from about 90.degree. to 120.degree. C. The prior art method such as treatment of lead sulfide with a large excess of aqueous acidic ferric chloride will give elemental sulfur but this ferric chloride medium is more corrosive in nature thus necessitating the use of more expensive equipment and, in addition, is not as selective in the chlorination of lead only, more impurity metals going into solution which will come over in the filtration step along with the soluble lead chloride. The halogenation of the lead sulfide as here proposed is effected at a temperature of from 90.degree. to about 120.degree. C. by treating said lead sulfide with a halogenating compound such as chlorine, bromine, fluorine, etc., in a dry gaseous state. The halogenation of the lead sulfide with the aforementioned halogen gas will result in the formation of a lead halide such as lead chloride, lead bromide, or lead fluoride with the attendant formation of elemental sulfur.
In the next step of the improved hydrometallurgical process the halogenated mixture is then subjected to a water wash wherein the soluble halides of the impurity metals which are present in the feedstock will be removed prior to subjecting the mixture to a brine leaching operation and thereby facilitate the recovery of metallic lead in a purer form. The water washing of the halogenated mixture will remove such soluble metal chlorides as ferric chloride, copper chloride, zinc chloride, cadmium chloride, etc., which had only slightly been halogenated, whereby the lead which is eventually recovered will be in a purer form than that which has heretofore been obtained. The water washing of the halogenated mixture may be effected over a relatively wide range of temperatures such as from about 5.degree. to about 95.degree. C., the amount of water which is utilized for the washing step varying according to the method of washing the halogenated mixture. The wash water is then separated from the solid halogenated mixture and charged to a treatment step whereby the wash water may be treated for discharge or may also, if so desired, be treated for the recovery of the metallic impurities which have been removed and separated from the halogenated lead mixture. The washed solids are then leached by the addition of a brine solution at an elevated temperature usually in the range of from about 80.degree. to about 120.degree. C., said brine solution usually comprising an aqueous sodium chloride solution containing from about 20 to about 35% by weight of sodium chloride. During the brine leaching step, the pH of the solution is maintained in the range of from about 4 to about 8 by the addition of acidic or caustic solution such as the hydroxides and oxides of Group IA of the Periodic Table including sodium hydroxide, potassium hydroxide, lithium hydroxide, Group IIA of calcium, magnesium oxide or halogen acids such as hydrochloric acid, hydrobromic acid, etc. By controlling the pH of the brine leaching solution in the aforesaid range, other metallic impurities which are present in the solution such as copper, silver, zinc, cadmium, antimony and possibly iron along with some unreacted sulfides will reprecipitate from the solution either by hydrolysis or by reaction to form insoluble sulfides under the conditions of the solution. The leaching of the mixture is effected for a period of time which may range from about 0.25 to about 2 hours or more in duration, the residence time being that which is sufficient to dissolve the lead halide.
Upon completion of the leaching step, the solution is filtered while maintaining the temperature of the solution at an elevated range of from about 80.degree. to about 120.degree. C. whereby the lead halide is maintained in a soluble form. The filtrate which contains the soluble lead halide is then passed to a crystallization zone wherein the soluble lead halide is crystallized due to a temperature drop, the solubility of the lead halide decreasing as the temperature decreases.
The thus crystallized lead halide is then recovered and, in the preferred embodiment of the invention, is dried to remove any trace of water which may still be present, the drying may be effected, if so desired, by placing the lead halide in an oven and subjecting the lead halide to a temperature of about 100.degree. C. in an atmosphere of air for a period ranging from about 0.1 to about 4 hours or more, the duration of the drying period being that which is sufficient to remove all traces of water. Following the drying of the lead halide, it is then placed in an appropriate apparatus and subjected to a temperature sufficient to melt said halide until it assumes a molten form. This temperature may range from about 380.degree. C. which is sufficient to melt lead bromide up to about 875.degree. C. which is sufficient to melt lead fluoride. The lead halide in molten form is then admixed with a salt of a metal selected from the group consisting of alkali metals and alkaline earth metals. Examples of these salts of metals of Groups IA and IIA of the Periodic Table will include lithium chloride, sodium chloride, potassium chloride, rubidium chloride, cesium chloride, beryllium chloride, magnesium chloride, calcium chloride, strontium chloride, barium chloride, lithium bromide, sodium bromide, potassium bromide, rubidium bromide, cesium bromide, beryllium bromide, magnesium bromide, calcium bromide, strontium bromide, barium bromide, lithium fluoride, sodium fluoride, potassium fluoride, rubidium fluoride, cesium fluoride, beryllium fluoride, magnesium fluoride, calcium fluoride, strontium fluoride, barium fluoride, etc., in a fused salt bath. In the preferred embodiment, the salt of a metal of Groups IA or IIA of the Periodic Table will be comparable in the halide content to the lead halide which is to undergo electrolysis, that is, if the lead halide is lead chloride, the solid salt will comprise a chloride such as sodium chloride, potassium chloride, lithium chloride, calcium chloride, etc. In general, the salt of the metal of Groups IA or IIA of the Periodic Table will be present in the fused salt mixture in an amount in the range of from about 20% to about 40% by weight of the mixture. It is also contemplated within the scope of this invention that the lead halide will undergo electrolysis in the presence of a mixture of at least two salts of the metals of Groups IA and IIA of the Periodic Table, examples of these mixtures comprising a sodium chloride-lithium chloride mixture, a potassium chloride-lithium chloride mixture, a magnesium chloride-calcium chloride mixture, a lithium bromide-potassium bromide mixture, etc. In the fused salt bath the mixture of salts will be subjected to electrolysis utilizing a sufficient voltage to effect said electrolysis whereby metallic lead is deposited as a liquid which can be removed from the fused salt. The lead may be removed continuously or batchwise.